United States Patent [19]
Lowenhaupt et al.
[11] Patent Number:
[45] Date of Patent:
4,541,868
Sep. 17, 1985
[54] RECOVERY OF NICKEL AND COBALT BY
CONTROLLED SULFURIC ACID LEACHING
[75] Inventors: E. Harris Lowenhaupt, Gasquet,
Calif.; John E. Litz, Lakewood;
Dennis L. Hoe, Broomfield, both of
Colo.
[73] Assignee: California Nickel Corporation,
Crescent City, Calif.
[21] Appl. No.: 516,235
[22] Filed: Jul. 22, 1983
[51] Int. Cl.4 COlG 53/00; COlG 55/00;
C22B 3/00
[52] U.S. Cl. 75/101 R; 75/108;
75/115; 75/119; 423/140; 423/141; 423/146;
423/150
[58] Field of Search 423/123, 140, 141, 146,
423/150,155; 75/101 R, 115, 119, 108
3,804,613 4/1974 Zundel et aI. 75/101 R
3,809,549 5/1974 Opratko 75/101 R
3,991,159 11/1976 Queneau et aI. 423/150
4,012,484 3/1977 Lussiez 423/53
4,044,096 8/1977 Queneau et aI. 423/150
4,065,542 12/1977 Subramanian et aI. 423/35
4,097,575 6/1978 Chou et aI. 423/150
4,098,870 7/1978 Fekete et aI. 423/123
4,195,065 3/1980 Duyvesteyn 423/150
4,410,498 10/1983 Hatch et aI. 423/150
OTHER PUBLICAnONS
Boldt, Jr., Joseph R., The Winning ofNickel, Longmans
Canada Limited, 1967, pp. 439-440.
"Freeport Nickel's Moa Bay Puts Cuba Among Ranking
Ni-Producing Nations", Engineering and Mining
Journal, vol. 160, No. 12, Dec. 1959, pp. 84-92.
Primary Examiner-John Doll
Assistant Examiner-Robert L. Stoll
Attorney, Agent, or Firm-Sheridan, Ross & McIntosh
11 Claims, No Drawings
According to the present invention, improved dissolution
of nickel and cobalt and thus improved recovery of
those desired metal values is achieved by modifying the
ore recovery processes wherein sulfuric acid leaching at
elevated temperatures is used to dissolve the nickel and
cobalt. In particular, according to the present invention,
processes are provided wherein the sulfuric acid and
ore are contacted at substantially ambient temperature
prior to subsequent heating to attain the elevated temperatures
of the sulfuric acid leach. Practice of the present
invention has been found to result in improved
metal value recovery.
[56] References Cited
U.S. PATENT DOCUMENTS
2,842,427 7/1958 Reynaud et aI. 23/183
2,872,306 2/1959 Morrow 75/101
2,971,836 2/1961 Hall 75/119
3,082,080 3/1963 Simons 75/115
3,093,559 6/1963 White 204/123
3,293,027 12/1966 Mackiw et aI 75/119
3,333,924 8/1967 Hazen et aI. 23/165
3,365,341 1/1968 Fitzhugh, Jr. et aI. 75/119
3,466,144 9/1969 Kay 23/183
3,473,920 10/1969 Fitzhugh, Jr. et aI. 75/109
3,720,749 3/1973 Taylor et aI. 423/141
3,737,307 6/1973 Fitzhugh, Jr. et aI. 75/109
3,761,566 9/1970 Michal 423/141
3,773,891 11/1973 O'Neill 423/139
3,793,432 2/1974 Weston 423/143
[57] ABSTRACT
4,541,868
1 2
RECOVERY OF NICKEL AND COBALT BY
CONTROLLED SULFURIC ACID LEACHING
DETAILED DESCRIPTION OF THE
INVENTION
EXAMPLE
A series of tests were performed comparing the effect
on nickel and cobalt dissolution by sulfuric acid leach-
The present invention provides for improved meth-
5 ods of recovering the metal values, i.e. nickel and cobalt,
from laterite ores wherein the ore is leached with
sulfuric acid at elevated temperatures to dissolve the
desired metal values and thereafter the metal values are
recovered from the leach liquor. In particular, the pro-
JO cess of the invention provides for the addition of the
sulfuric acid to the ore, i.e. contacting the ore with
sulfuric acid, prior to heating or otherwise attaining the
desired elevated temperature for leaching.
As described hereinabove, a number of different processes
utilizing sulfuric acid leaching are known for
recovery of nickel and cobalt from ores such as lateritic
ores. The benefits of the present invention can be
achieved in conjunction with virtually any process
wherein the sulfuric acid leach is at elevated temperatures.
In general, feed material to processes in which the
present invention can be applied is a laterite ore containing
some or all of the following major constituents:
peridotite, saprolite, hematite, magnetite, geothite, garnierite,
maghemite, and aluminum oxides. Typically,
the laterite ore is slurried and sized by conventional
methods. Some or all of the ore is contacted at elevated
temperatures and pressures with sulfuric acid, typically
at temperatures of from about 200· to about 270· C. and
more typically at about 240· c., and pressures of from
400 to 700 psig. In general, the leach will continue for a
time period sufficient to solubilize or leach the nickel
and cobalt present into solution. The specific parameters
of the various leach processes described in U.S. Pat.
Nos. 4,044,096, 4,195,065 and 3,466,144 are incorporated
herein by reference.
The leach slurry which comprises the leach residue
and a leach liquor is then separated by conventional
liquid/solid separation means such as by countercurrent
decantation. The liquor is then advantageously concentrated
by evaporator before being neutralized. Typical
neutralizing agents are magnesium oxide or hydroxide
and/or magnesium-rich ore or ore fractions. Following
neutralization the liquor is again separated from any
solids present and may be thickened and/or concentrated
by evaporator before passing to conventional
metal recovery, such as contact with H2S and/or H2 to
precipitate Ni and Co, or by adding a neutralizing agent
to precipitate hydroxides of Co and/or Ni.
As will be understood by those skilled in the art, the
process of the present invention wherein the H2S04 and
ore are contacted at lower than conventional temperatures
may be applied to any of the known processes
described generally hereinabove to achieve enhanced
55 Co and Ni dissolution into H2S04 leach liquor. In the
preferred embodiment of the present invention the ore
and H2S04 are contacted at ambient conditions, i.e.
room temperature frequently from about 20· to about
30· C. and at atmospheric pressure. However, improved
60 dissolution may be achieved by contacting at virtually
any temperature from about ambient to below the temperature
of the subsequent leach step.
The following example is provided by way ofillustration
and not by way of limitation.
SUMMARY OF THE INVENTION
BRIEF DESCRIPTION OF THE PRIOR ART
FIELD OF INVENTION
This invention relates to the recovery of nickel and
cobalt from lateritic ores, and, in particular, to a method
of sulfuric acid leaching which maximizes the solubility
of the desired nickel and cobalt values.
A wide variety of processes are known for the recovery
of nickel and/or cobalt from various nickel-bearing
ores including laterite and serpentine ores. Basic to one 15
class of processes is the solubilizing of the nickel and/or
cobalt by sulfuric acid leaching followed by neutralization.
Since the sulfuric acid leach is central to the extraction,
a number of different process parameters and
conditions have been developed in order to maximize 20
dissolution of the desired metals during the leach.
U.S. Pat. No. 3,991,159 is representative of prior art
processes wherein sulfuric acid leaching is carried out
at elevated temperatures, e.g. 200 to about 300· c., and
elevated pressures, e.g. 225 to 1750 psig. Other exam- 25
pIes of such sulfuric acid leaching are disclosed in U.S.
Pat. Nos. 4,044,096, 4,195,065, and 3,466,144.
In each of these prior art references, it is taught that
the sulfuric acid should be added to an already heated
ore slurry. For example, in U.S. Pat. No. 4,195,065, the 30
ore slurry is preheated to a temperature of 230·-300· C.
before the sulfuric acid is added and a leach at that
temperature is performed. In U.S. Pat. No. 3,466,144
the slurry is preheated to 400·-500· F. before acid addition.
Similarly, in U.S. Pat. No. 4,044,096 the acid addi- 35
tion is to a preheated slurry in or at the entrance of the
autoclave. In fact, adding the sulfuric acid to a preheated
slurry has been considered the preferred method
for as long as these processes have been known and 40
used. Many existing U.S. patents described and are
based upon the process developed in connection with
operations at Moa Bay, Cuba. These operations are
described in detail in the article entitled "Freeport
Nickel's Moa Bay Puts Cuba Among Ranking 'Ni-Pro- 45
ducing' Nations", Engineering and Mining Journal, Vol.
160, No. 12, December 1959, pp. 84-92. The article
specifically teaches that the best results are obtained
when sulfuric acid is added to the slurry after the leach
reaction temperature has been obtained. 50
It has now been discovered that cold addition of
H2S04 prior to attaining the leach reaction temperature
advantageously affects Ni and Co dissolution and recovery.
According to the present invention, improved dissolution
of nickel and cobalt and thus improved recovery
of those desired metal values is achieved by modifying
the ore recovery processes wherein sulfuric acid leaching
at elevated temperatures is used to dissolve the
nickel and cobalt. In particular, according to the present
invention, processes are provided wherein the sulfuric
acid and ore are contacted at substantially ambient temperature
prior to subsequent heating to attain the ele- 65
vated temperatures of the sulfuric acid leach. Practice
of the present invention has been found to result in
improved metal value recovery.
Sulfuric add Residue Dissolution
Test No. addition (IblTon) %Ni % Co %Ni % Co
IA 675l! 0.16 0.017 85 89 10
B 6752/ 0.085 0.012 92 92
2A 700 1/ 0.105 0.014 90.3 90.0
B 7002/ 0.087 0.006 91.3 95.1
3A 915 1/ 0.099 0.009 92.0 94.1
B 91021 0.047 0.005 96.3 96.2
I/lnjcctcd into pulp at Icuch l,:onditions. 15
~/Added to pulp at ambient temperature and pressures prior 10 advancing into IC'1Ch
autoclave.
4,541,868
35
3
ing when H2S04 was added before and after reaching
the elevated leach temperature. Leaching in each instance
was for 60 minutes at 240· C. and 550 psig. The
leach residue was analyzed and results obtained are
tabulated in Table 1.
TABLE I
Although the foregoing invention has been described
in some detail by way of illustration and example for
purposes of clarity and understanding, it will be obvious
that certain changes and modifications may be practiced
within the scope of the invention, as limited only by the
scope of the appended claims.
What is claimed is:
1. A method of recovering nickel and cobalt from
lateritic ores by sulfuric acid leaching comprising:
(a) mixing the sulfuric acid for said leaching with said
ore at a temperature from about ambient temperature
to below the leach temperature of step (b);
(b) leaching said nickel and cobalt from said ore at a
temperature above about 200· C. to form a nickeland
cobalt-containing pregnant leach liquor and a
leached residue with no sulfuric acid addition at
said leach temperature;
(c) separating said pregnant leach liquor from said
residue; and
(d) recovering said nickel and cobalt from said pregnant
leach liquor.
2. A method according to claim 1 wherein said ore is
a laterite ore fraction of less than - I inch in size.
3. A method according to claim 11 further comprising
mixing said sulfuric acid and said ore at at temperature
below about 160· C.
4. A method according to claim 3 further comprising
heating the mixture of ore and sulfuric acid to a temper-
4
ature of from about 200· C. to about 270· C. whereby
nickel and cobalt are dissolved into a pregnant liquor.
5. A method according to claim 3 further comprising
contacting said pregnant leach liquor with H2S to preS
cipitate NiS.
6. A method according to claim 3 further comprising
contacting pregnant leach liquor with H2 to reduce any
cobalt present to its elem~ntal state.
7. A method according to claim 3 further comprising
contacting said pregnant leach liquor with a base to
precipitate the hydroxides of any nickel and cobalt
present.
8. In a method of recovering nickel and cobalt from
lateritic ores comprising contacting said ore with an
H2S04 leach liquor at a temperature of from about 200·
C. to about 270· C. to solubilize said nickel and cobalt
into said leach liquor and thereafter recovering said
nickel and cobalt from said leach liquor, the improvement
comprising adding substantially all the H2S04 of
20 said leach liquor to said ore at a temperature from about
ambient to below about 160· C. and thereafter heating
said mixture to a temperature of from about 2000 C. to
about 300· C.
9. The improvement of claim 8 wherein said adding is
25 at ambient temperature and atmospheric pressure.
10. In a method of enhancing the dissolution of nickel
and cobalt from nickel- and cobalt-containing lateritic
ore into a H2S04 leach liquor comprising adding the
H2S04 of said leach liquor to said ore at about ambient
30 temperature and atmospheric pressure and heating said
mixture to a temperature of from about 200· C. to about
300· C. for a time period sufficient to solubilize substantially
all of said nickel and cobalt with no further addition
of H2S04 after said heating.
11. In a method of enchancing the dissolution of
nickel and cobalt from nickel- and cobalt containing
lateritic ore into a sulfuric acid leach liquor wherein
said H2S04 leach liquor is contacted with said ore at a
temperature from about 200· C. to about 300· C. for a
40 time period sufficient to solubilize substantially all of
said nickel and cobalt, the improvement comprising
adding the H2S04 of said leach liquor to said ore at
about ambient temperature and atmospheric pressure so
that substantially all of the sulfuric acid is present dur-
45 ing the heating of the ore to said temperature.
* * * * *
50
55
60
65
atchK��:rtPA@swell
as for continuous operations, and to control the rate of
solids removal.
Returning to FIG. 2, a method for facilitating removal
of the metals from the pregnant leaching solution
withdrawn from the stilling well 19 will now be described.
Pregnant leaching solution is withdrawn from stilling
well 19 through line 42, and passes to the carbon adsorbing
device 43, ultimately passing to line 25.
Operatively connected to the device 43 via line 44 is
the storage tank 45, which is adapted to store strong
leaching solution therein. The invention is particularly
applicable to the leaching of gold and silver ores, in
which case the leaching liquid from source 26 comprises
a caustic cyanide solution, such as sodium cyanide,
and the solution in tank 45 is supplied from a
source 46 of hot caustic cyanide solution. Fresh carbon
is periodically added to the top of device 43 via line 47
and loaded carbon is withdrawn from the bottom
through line 44 and dumped into tank 45.
From tank 45, caustic and carbon are periodically
passed by pump 49 through valve 50 to metal removal
station 51. From metal removal station 51 carbon is
passed to carbon reactivation station 52 to supply fresh
carbon, through line 47, to device 43 as needed.
3
particles in a stable network. The slurry is passed to a
leach tank for counter-current flow treatment with a
leaching liquid. The pregnant solution is passed to a
station for metal recovery, while the leached slurry is
passed to wash tanks, with the residue passed to a tail- 5
ings pond, or like disposal area.
FIG. 2 schematically illustrates equipment that may
be utilizable for the practice of the novel aspects of the
process illustrated by the flow sheet of FIG. 1. The
particlized ore is passed by conveyor 10 or the like to IO
the slurrying tank 11, which may include a conventional
mixing means 12. A slurrying liquid is added to the tank
11 from line 13, the liquid in line 13 preferably comprising
water and/or the spent wash liquid from one or
more washing stages to be hereinafter described.
Also added to the slurry tank 11 is a flocculant from
source 14, and fibers from source 15. The flocculant and
fibers may be any suitable flocculant or fibers that are
capable of locking ore particles, including fines, in the
slurry in a stable network so that they may be subse- 20
quently subjected to the leaching treatment. Typical
flocculants comprise synthetic polymers of anionic,
cationic or nonionic types, and typical fibers comprise
cellulosic fibers, fiberglass fibers, ceramic fibers, and
combinations thereof.
The slurry from tank 11 is passed through line 17 to
the top of a substantially vertically elongated leaching
reactor vessel 18. The vessel may be of a type such as
disclosed in U.S. Pat. Nos. 4,061,193 or 4,174,997, having
a "stilling well" structure 19 at the top thereof 30
above the slurry introducing point 20 of a slurrry introduction
tube 21, and having a rotating liquid introducing
device 22 at the bottom thereof, preferably commonly
rotatable with slurry discharge structure 23.
The slurry flows continuously downwardly in vessel 35
18, and is ultimately discharged through line 24 at a
bottom portion of the vessel 18. Leaching liquid is introduced
through line 25-as from source 26-so that it
flows to the distributor 22, and then flows upwardly in
vessel 18-counter-currently to the slurry flow therein. 40
The fibers and flocculant in the ore slurry lock the ore
particles in a stable network so that leaching can occur
without channelling, and with a minimal pressure drop
over the vessel 18. Thus fines of 200 mesh, or smaller,
can be handled without substantial difficulty. Addition- 45
ally, ores having particle sizes up to i inch mesh, or
even larger, may be handled at the same time that the
fines are being handled.
The leached slurry withdrawn in line 24 passes to a
washing station. At the washing station, preferably a 50
single washer 27 is provided, although a first washer
(27) and a second washer 28 (or more) may be provided.
The vessels 27, 28 are substantially identical to the vessel
18, but generally of smaller size, and include a stilling
well arrangement 29, 29' at the top thereof, and a 55
rotating liquid introducing structure 30, 30' at the bottom
thereof. Where two vessels are utilized, clean wash
water enters vessel 28 through line 31, passes through
liquid distributor 30' counter-current to the slurry flow
in vessel 28, and spent wash liquid is withdrawn from 60
the stilling well 29' through line 32, to be used as feed
wash liquid for the first wash vessel 27. Spent wash
liquid in vessel 27 withdrawn from stilling well 29
through conduit 33 preferably is passed to the slurrying
tank 11, providing a liquid feed to line 13. The washed 65
slurry is withdrawn from second wash tank 28 through
line 34, and is passed to a tailings pond, or like disposal
site.
4,501,721
5
ling well 19-without the necessity for utilizing screens-
in line 42.
The pregnant solution in line 42 passes through the
carbon adsorber device 43, and then is recirculated
through line 25 to the device 22. Fresh carbon is sup- 5
plied to, and loaded carbon removed from, device 43 as
necessary.
Leached slurry is discharged by device 23 from the
bottom of the vessel 18 into line 24, and passes to the top
of washing stage 27. The slurry flows downwardly 10
through washing stage 27, and then may be fed to the
top of vessel 28, and ultimately passes through discharge
34 to a disposal site. Wash water introduced in
line 31 to vessel 28 flows counter-currently to the
slurry, and the spent wash liquid withdrawn from line 15
33 is pumped to the slurry in tank 11.
Utilizing the equipment illustrated in FIGS. 2 and 3,
it is easy to adapt the process to high temperature and
high pressures, thus providing versatility in the processes
which may be practiced, and in the available 20
leaching liquids. Also, the process can be closed to the
atmosphere thereby minimizing discharges of pollutants
into the atmosphere. Because little agitation of the
slurry is necessary (only that small amount provided in 25
tank 11), energy requirements are minimized, and the
process is readily adaptable to the handling of "soft"
particlized mineral materials, such as coal. Further, the
process can be operated in a continuous manner even
when a relatively high percentage of fines, including 30
small fines, are present, and may be practiced with particles
up to about ~ inch in diameter.
While the invention has been herein shown and described
in what is presently conceived to be the most
practical and preferred embodiment thereof, it will be 35
apparent to those of ordinary skill in the art that many
modifications may be made thereof within the scope of
the invention, which scope is to be accorded the broadest
interpretation of the appended claims so as to encompass
all equivalent methods and products. 40
What is claimed is:
1. A method of removing metal from a particlized
metal bearing ore, utilizing a treatment vessel, comprising
the steps of:
(a) mixing the particlized ore with a liquid to form a 45
liquid slurry, a flocculating material so as to lock
the particlized ore in a stable network in the slurry,
and fibers so as to facilitate locking of the particlized
ore in a stable network in the slurry;
(b) continuously passing the slurry downwardly in 50
the vessel;
(c) continuously passing a leaching liquid, capable of
leaching the metal to be removed from the metal
bearing ore, upwardly in the vessel, countercurrent
to the slurry passage, to remove metal from the ore 55
in the slurry;
(d) continuously removing the treatment slurry from
a bottom portion of the vessel;
6
(e) continuously removing pregnant leaching liquid,
with metal removed from the ore, from a top portion
of the vessel; and
(f) continuously washing the slurry removed in step
(d).
2. A method as recited in claim 1 wherein the flocculent
is selected from the group consisting essentially of
synthetic polymers of anionic, cationic, and nonionic
types.
3. A method as recited in claim 1 wherein the fibers
are selected from the group consisting essentially of
cellulosic fibers, fiberglass fibers, ceramic fibers, and
mixtures thereof.
4. A method as recited in claim 3 wherein the flocculent
is selected from the group consisting essentially of
synthetic polymers of anionic, cationic, and nonionic
types.
5. A method as recited in claim 4 wherein the fibers
comprise, by weight, between about 0.01 percent and 10
percent of the slurry.
6. A method as recited in claim 5 wherein the fibers
comprise, by weight, between about 0.05 and 0.75 percent
of the slurry.
7. A method as recited in claim 4 wherein the leaching
liquid in step (c) is a cyanide solution.
8. A method as recited in claim 1 wherein the fibers
comprise, by weight, between about 0.01 percent and 10
percent of the slurry.
9. A method as recited in claim 8 wherein the fibers
comprise, by weight, between about 0.05 and 0.75 percent
of the slurry.
10. A method as recited in claim 1 wherein step (f) is
practiced utilizing a washing vessel, and by passing
wash liquid upwardly in the vessel counter-current to
slurry moving downwardly in the vessel, with spent
wash liquid removed from a top portion of the vessel
and washed slurry removed from a bottom portion of
the vessel.
11. A method as recited in claim 1 consisting essentially
of said steps (a)-(f), so that carbon addition to the
slurry is not practiced.
12. A method as recited in claim 10 wherein said
washing vessel comprises a first vessel, and wherein
step (f) is practiced utilizing a second vessel substantially
identical to the first vessel, the slurry outlet from
the first vessel being connected to the slurry inlet to the
second vessel, and the wash liquid inlet to the first vessel
being connected to the wash liquid outlet from the
second vessel; and wherein the wash liquid outlet from
the first vessel is used as a liquid for practicing step (a).
13. A method as recited in claim 1 wherein the leaching
liquid in step (c) is a cyanide solution.
14. A method as recited in claim 13 comprising the
further step of: (g) continuously passing pregnant liquid
withdrawn in step (e) through a carbon adsorbing device,
and reintroducing the liquid into the vessel as
leaching liquid for step (c).
* * * * *
60
65
imes zw�^RmPA@sf";mso-fareast-font-family: HiddenHorzOCR'>Co. Ni rougher (I)
(2)
Co. Ni 1st cleaner
Co, Ni 2nd cleaner
Co, Ni 3rd cleaner
Stage
Equipment
Speed (rpm)
Ca(OHh
0.10
0.05
0.05
0.05
Rougher
lOOOgD-1
1800
NaCN
0.05
0.025
0.025
0.025
0.05
0.025
0.025
0.D25
0.01
0.5
0.05
0.05 0.2
0.01
0.01
0.01
Co, Ni 1st cleaner
500 g D-I
1500
MIBC4 Grind Cond Froth pH
I 2 9.5
I 2
I I
I I 9.5
5 8.5
I 4 8.5
2 4
I 4 8.0
I 3
I 2
Remaining cleaners
250 g D-I
1200
1Ethyl isopropyl thionocarbamate
2Ammonium diisopropyl dithiophosphate
3Sodium isopropyl xanthate
-4Methyl isobutyl carbinol
sPolUSium amyl xanthate
Example IV-Table 5 summarizes the results obtained
from cycle testing according to Examples I, II and III.
As much as 91% of the copper, 85% of the lead and
92% of the cobalt and nickel values were recovered in 25
their respective concentrates. Cycle tests were not con·
ducted on Samples 1 and 4. A primary grind of 60 to
70% passing 200 mesh was employed. Thickening and
filtration rates of the products were judged adequate to
good.
effecting flotation of the copper and separating a
copper rougher concentrate from a copper rougher
tailing product;
regrinding the copper rougher concentrate to liberate
lead and cobalt-nickel minerals and conditioning
the reground concentrate with S02;
cleaning the reground conditioned rougher concentrate
and separating a first copper cleaner concentrate
from a first copper cleaner tailing product;
TABLE 5
Weight Assays, % Distribution, %
Product % Cu Pb Co Ni Cu Pb Co Ni
Sample No.2
Cu conc 2.51 28.6 4.68 0.19 0.27 89.0 11.6 3.3 3.0
Pb conc 1.01 0.84 79.2 0.14 0.18 1.0 78.9 1.0 0.8
Co-Ni COnc 3.24 1.16 1.05 3.80 5.85 4.7 3.4 86.1 82.5
Head (calc) 0.81 1.01 0.143 0.23
Sample No.3
Cu conc 3.25 27.6 4.75 0.23 0.32 89.0 9.1 4.2 4.0
Pb conc 1.70 0.30 84.8 0.11 0.15 0.5 85.0 1.1 1.0
Co-Ni conc 5.38 1.17 0.91 2.70 3.85 6.2 2.9 81.2 80.4
Head (calc) 1.01 1.69 0.179 0.26
Sample No.5
Cu conc 6.84 31.2 2.32 0.25 0.32 90.9 10.5 3.2 3.2
Pb conc 1.64 0.56 78.6 0.28 0.38 0.4 85.1 0.9 0.9
Co-Ni conc 5.95 2.59 0.62 8.30 10.6 6.5 2.4 92.4 91.7
Head (calc) 2.35 1.51 0.53 0.69
50 routing at least the copper rougher tailing product
directly to the lead flotation circuit wherein a lead
concentrate is separated from a lead tailing product;
routing the lead tailing product from the lead flotation
circuit to a cobalt-nickel flotation circuit
wherein a cobalt-nickel concentrate is separated
from a cobalt-nickel tailing product; and
recovering the copper, lead and cobalt-nickel concentrates
from their respective flotation circuits.
2. The invention of claim 1, wherein the copper
rougher tailing product and first copper cleaner tailing
product are combined and routed to the lead flotation
circuit.
3. The invention of claim 1, wherein flotation of cop65
per is effected in the absence of pH modifiers other than
sulfur dioxide or sulfurous acid.
4. The invention of claim 1, wherein the primary
grind pulp is conditioned by addition of S02 in an
What is claimed is:
1. In a sequential flotation process for the separation
of components of a mineral mixture of the type wherein
a primary grind ore pulp is routed sequentially through
a series of flotation circuits having successive separation 55
and concentration stages for separating and concentrating
one of the mineral components, the improvement
comprising:
grinding a sulfide ore comprising a mixture of copper,
lead and cobalt-nickel sulfide minerals in a carbon- 60
ate matrix to provide a primary grind flotation
pulp;
conditioning the pulp with S02 under intense aeration
to depress lead and cobalt-nickel and promote copper;
routing the conditioned pulp to a copper flotation
circuit having a roughing stage and at least one
cleaning stage;
10
8. The invention of claim 1, wherein the sulfide ore is
a Missouri lead belt ore.
9. The invention of claim 1, wherein the sulfide ore is
a vibumam trend ore body of the new lead belt.
10. The invention of claim 1, wherein the sulfide ore
is located within a Mississippi Valley-type deposit.
11. The invention of claim 1, wherein the flotation of
copper is effected at an acidic pH of about 6.5 to 6.8.
12. The invention of claim 11, wherein a collector
10 highly preferential for copper in an acidic medium is
employed for copper flotation.
13. The invention ofclaim 11, wherein the collector is
ethyl isopropyl thionocarbamate.
* * * * *
4,460,459
9
amount of from about 1 to about SIbs. 802 per ton of
pulp.
S. The invention of claim 1, wherein the primary
grind pulp is intensely aerated by injection of natural air 5
into the pulp at a rate of about 3 to 5 cu ft/min.
6. The invention of claim 1, wherein lead is separated
by flotation after depression of other sulfides present
with a cyanide.
7. The invention of claim 1, wherein cobalt/nickel is
separated by flotation after activation with copper sulfate.
15
20
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30
35
40
45
50
55
60
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t: �Z{a;PA@sation:none;mso-layout-grid-align:none;text-autospace:none'>4,402,919
Sep. 6, 1983
• • '. •• I ,.. _ '. _ "- • •
. -
1
4,402,919
2
SUMMARY OF THE INVENTION
DESCRIPTION OF THE PREFERRED
EMBODIMENT
A feed material comprising an ore containing aluminum,
phosphorus and other values including uranium is
treated to obtain a select fraction having a particle size
less than about 150 U.S. Standard mesh and preferably
less than 200 U.S. Standard mesh to provide a concentrate
fraction. The concentrate fraction contains valuable
quantities of uranium and other elements such as
aluminum and phosphorus. The remainder of the feed
material predominately comprises quartz and is discarded.
The treatment by which the concentrate fraction
is obtained can include crushing, scrubbing, grinding
or milling of the ore to provide a particulate capable
of being sized. The particulate is sized by screening or
any other suitable means. The particular apparatus employed
can comprise any commercially available equipment
capable of producing the concentrate fraction.
fraction of large particles size quartz sand and upgrades
the unlnium, aluminum and phosphorus content of the
remainder. Acid consumption still is substantially
higher than desirable. For example, 1600 to 2000
5 pounds of 93 to 98 percent sulfuric acid are required to
dissolve 2000 pounds of concentrate. It is known that
calcining the concentrate before dissolution will reduce
acid consumption. However, acid consumption remains
at about 600 pounds of93 to 98 percent sulfuric acid per
2000 pounds of original concentrate.
It is desirable to provide a process that will permit
regeneration of a portion of the acid that is consumed to
solubilize the ore in which the uranium is present.
The surprising discovery now has been made that
uranium can be dissolved from an ore comprising aluminum,
phosphorus, uranium and other values by a procedure
which reduces the quantity of acid consumed to
effect the dissolution by over one half the quantity presently
consumed in the best prior art process. The reduction
in acid consumption is effected by regeneration of
a substantial portion of the acid consumed to solubilize
the ore.
In practice, the ore is contacted with a mineral acid to
solubilize at least a portion of the acid soluble constituents
including any uranium contained in the ore. The
ore can be physically concentrated or otherwise treated
30 such as by calcination prior to contacting the mineral
acid. The solubilization results in the formation ofa
spent acid solution containing dissolved uranium, aluminum,
phosphorus and other values together with any
undissolved solids. The spent acid solution then is
heated to a predetermined elevated temperature while
maintaining at least the autogenic pressure of the solution
to effect a precipitation of aluminum phosphate
from the solution. The precipitation results in the regeneration
of a substantial portion of the mineral acid consumed
to solubilize the ore. The uranium values then
can be recovered from the remaining solution by any
known techniques. The uranium depleted solution comprising
regenerated acid then is recycled to contact
fresh ore to solubilize additional uranium values.
Alternatively, the uranium can be recovered from the
spent acid solution prior to acid regeneration.
An additional benefit of the process is the production
of a high quality aluminum phosphate by-product.
BACKGROUND OF THE INVENTION
PROCESS FOR THE REGENERATION OF
MINERAL ACIDS USED TO SOLUBILIZE
PHOSPHATE ORES
1. Field of the Invention
This invention relates to a process for the regeneration
of mineral acids used to solubilize phosphate ores
which thereby permits recovery of uranium and other 10
valuable minerals from the ore.
2. Description of the Prior Art
It is well known in the phospheric acid technology
that phosphate ore can be treated with a mineral acid to
convert the phosphate into a soluble form, either as 15
phosphate fertilizers, phosphoric acid or phosphoric
acid compositions which can be processed into phosphate
chemicals. The solubilization process also is
known to dissolve impurities in the ore such as uranium
and vanadium which then can be separately recovered 20
from the resultant solution. One of the largest economic
expenses of the process is the cost of the mineral acid
that is consumed during the solubilization. The quantity
of mineral acid required to effect the solubilization is
directly related to the quantity of acid soluble materials 25
present in the ore. Most of the acid soluble materials are
dissolved in the process of solubilizing the phosphate
values. No simple method is known in the prior art to
regenerate the acid used to convert the phosphates into
a soluble form.
Large phosphate ore fields are known to exist in
Florida and in other areas of the United States. For
economic reasons, only the phosphate ores containing a
high ratio of phosphate to other acid soluble materials
are considered commercially recoverable. The high 35
quality commercially recoverable ores of the Florida
fields have been found to contain limited quantities of
uranium. The overburden on the high quality phosphate
ore comprises material referred to as "leached zone
material" which consists largely of sand containing 40
components of aluminum, phosphorus, iron and other
values together with clays. The leached zone material
has been formed by natural weathering or leaching of
the phosphate ore field. The low phosphate content of
this leached ore presently makes its utilization unattrac- 45
tive for the production of phosphates because of the
large quantity of mineral acid required to solubilize the
ore. However, this leached ore has been found to contain
uranium in concentrations significantly greater than
in the higher quality phosphate ore that is considered 50
commercially recoverable.
The major problem preventing the recovery of the
uranium in the Florida leached zone material and from
other phosphate ore fields is one of economics. A large
quantity of acid is required to effect dissolution of the 55
uranium present in these ores. The high acid requirement
is due to the fact that the aluminum, phosphorus
and other acid soluble values also must be dissolved to
solubilize the uranium. Further, no effective method of
physically concentrating the minerals to produce a sig- 60
nificantly higher quality concentrate for treatment has
been found.
Presently, the best known concentrating procedure
produces a concentrate of the uranium and other phosphate
minerals by scrubbing and sizing the raw ore to 65
obtain a select fraction which then is dissolved with a
mineral acid. This procedure rejects from about 60 to
about 75 percent of the ore, by weight, as a coarse
~ ~ • • "I _ -_ .. • ' .... •
. . . - . .' . ~ ..
4,402,919
3
The concentrate fraction then is admixed with a sufficient
quantity of a leach solution comprising a mineral
acid to effect solubilization of a substantial portion of
the concentrate fraction and at least a portion of the
uranium present in said concentrate fraction. The min- 5
eral acid can comprise, for example, sulfuric acid, phosphoric
acid and the like. In a typical reaction, alSO
mesh size fraction in aqueous slurry form, having a
solids content in the range of from about 30 percent to
about 60 percent, is reacted with the sulfuric acid at 10
temperatures in a range between about ambient temperature
to above the boiling temperature of the leach
solution and preferably from about 60° C. to about 90°
C. For temperatures above the boiling temperature of
the leach solution, the solubilization is effected under a 15
pressure at least equal to the autogenic pressure of the
heated solution.
Preferably, the solubilization is carried out for a period
of time ranging between 0.2 and about 15 hours and
more particularly, for a period of from about 30 minutes 20
to about 60 minutes, although the length of time may be
varied considerably depending upon other variables in
the reaction conditions. The interdependence of variables
makes for vast differences in the specific conditions
employed as to each variation. In general, it may 25
be stated the higher the percent acid acidulation used,
the shorter the time required. Thus, for example, if
about 70 percent acidulation is used, that is, about 106.5
pounds of 96 percent sulfuric acid per 100 pounds of
ore, only about 15 minutes is required to acomplish the 30
digestion, while at about 45 percent acidulation, about 6
hours digestion is necessary to give good recovery of
the desired constituents. Depending upon the analysis
of the particular ore processed, between about 30 percent
and 105 percent acidulation is desired. This corre- 35
sponds to the addition of between about 29 pounds and
about 150 pounds of sulfuric acid per hundred pounds of
ore processed. Preferably, about 70 percent acidulation
is used. The percent acidulation referred to in this description
is calculated on the basis of the reaction of 40
sulfuric acid with all of the aluminum, calcium and iron,
or other significant cationic constituents present in the
ore. In other words, 100 percent acidulation would be
the addition of that amount of sulfuric acid required to
completely react with these components. After the 45
solubilization, the aqueous solution of reaction products,
sometimes referred to as "spent acid solution," is
separated from the insolubles, such as quartz and clay.
The substantially solids-free aqueous solution of reaction
products is introduced into a reaction zone wherein 50
the solution is heated to a temperature above 100° C.
while maintaining the solution at a pressure level at least
equal to the autogenic pressure of the solution to effect
a precipitation of the leached phosphorus values as
crystalline aluminum phosphate. Preferably, the aque- 55
ous solution is heated to a temperature level in the range
offrom about 150° C. to about 200° C. and most preferably
a temperature in the range of from about 180° C. to
about 200° C. Temperatures above 200° C. can be employed
to effect the precipitation of the leached phos- 60
phorus values, however, the precipitation reaction is
essentially complete at about 200° C.
The present inventors have found that when the
leached phosphorus values are precipitated within the
aqueous solution, in the described manner, that a por- 65
tion of the mineral acid is regenerated. This is evidenced
by a significant drop in the pH level of the aqueous
solution of reaction products as the aluminum phos-
4
phate precipitate is formed. The aqueous slurry produced
at a result of the precipitation of the AIP04 also
contains other values, including uranium, that were
dissolved during solubilization of the ore. These additional
elements remain in the solution and generally do
not precipitate with the aluminum phosphate.
While the precise mechanism of the chemical reaction
involved in regeneration of the mineral acid presently
in unknown, the inventors presently believe that
the major portion of the aluminum and phosphorus
contained in the aqueous solution of reaction products,
resulting from solubilization of the ore, is in the form of
AIH2P04+2. It is believed that the mineral acid is regenerated
according to the following equation:
AIHZP04+Z+mineral acid
anion~AIP04ppl+ mineral acid
More particularly, when sulfuric acid is employed to
solubilize the ore, the acid is believed to be regenerated
according to the following equation:
An analysis of the precipitate employing x-ray diffraction
indicates that the precipitate comprises berlinite, an
anhydrous aluminum phosphate. Further, chemical
analysis of the precipitate indicates that it contains no
detectable quantity of uranium and no significant quantity
of any of the other solubilized mineral values present
in the aqueous solution of reaction products, the
precipitate being found to have a purity in excess of 99
percent aluminum phosphate. Thus, the process of this
invention also produces a high quality by-product that
has a significantly higher P20S content than, for example,
apatite, which is considered a high quality source of
phosphorus.
The precipitated aluminum phosphate can be separated
from the aqueous slurry of the same by filtration,
centrifugation gravity settling or the like. The particular
apparatus employed to effect the separation can
comprise any of that which commercially is available.
In one particular embodiment in which the mineral
acid comprises sulfuric acid, if an attempt is made to
precipitate the aluminum and phosphorus from the
aqueous solution of reaction products at a temperature
below about 100° C., a precipitate will form. However,
the precipitate is alunogen (Ab(S04h.18H20) and no
mineral acid is regenerated. When sulfuric acid comprises
the mineral acid, it also has been observed that
any calcium sulfate which may tend to precipitate from
the aqueous solution of reaction products after formation
of such solution should be permitted to form. The
precipitated calcium sulfate then should be separated
from the aqueous solution before introduction of the
now substantially solids-free aqueous solution into the
reaction zone to precipitate the aluminum phosphate.
Otherwise, a mixed calcium aluminum sulfate is found
to precipitate instead of aluminum phosphate and no
sulfuric acid is regenerated.
The presence of calcium in the aqueous solution
when mineral acids other than aulfuric acid are employed
to effect the solubilization of the ore has no
apparent effect upon the precipitation of the aluminum
and phosphorus values as aluminum phosphate. The
filtrate remaining after separation of the aluminum
phosphate, which contains dissolved uranium and other
4,402,919
EXAMPLE II
5
elements, can be treated by any known method to recover
the uranium and any other desired elements.
The uranium can be separated from the filtrate by, for
example, solvent extraction techniques whereby the
uranium values are transferred from the aqueous filtrate 5
to an organic solvent extractant. The extracted uranium
then is separated from the organic solvent by, for example,
contact with an alkaline stripping agent. Various
processes for solvent extraction of tranium and other
values from aqueous acidic solutions are disclosed in, 10
for example, U.S. Pat. Nos. 3,700,415, 3,711,591 and
3,836,476, the disclosures of which are incorporated
herein by reference. It is to be understood that the
method for separating the uranium or any other values
from the aqueous solution is not to be limited to solvent 15
extraction processes but that any method known by
individuals skilled in the art may be employed.
The practice of the process of the present invention
results in the regeneration of over 50 percent of the acid
employed to solubilize the ore. Often, the present pro- 20
cess effects regeneration of over two thirds of the mineral
acid originally employed to solubilize the ore. Such
regeneration capability permits applicants to recover
uranium present in low phosphate content ores in an
economical manner while also providing a high purity 25
by-product of aluminum phosphate which can be used
as a feed stock for production of aluminum and phosphorus
chemicals.
To further illustrate the process of the present invention,
and not by way of limitation, the following exam- 30
pIes are provided.
EXAMPLE I
A representative sample of an aqueous solution of
reaction products resulting from sulfuric acid leaching 35
of a minus 150 mesh fraction of Florida leached zone
material is introduced into a reaction zone comprising a
Parr autoclave having an acid resistant liner. The solution
is formed by contacting 1600 lbs. of 96 percent
H2S04 with one ton of uncalcined leached zone mate- 40
rial. The aqueous solution is analyzed and is found to
contain 55.3 gil Ah03, 30 gil P20S, 0.11 gil U30g and
have a pH of about 0.5. The aqueous solution is heated
in the reaction zone to a temperature of about 200· C.
while maintaining the autogenic pressure of the aqueous 45
solution. The solution· is maintained at the elevated
temperature for about 5 minutes to effect precipitation
of crystalline aluminum phosphate in the solution to
form a slurry. The slurry is withdrawn from the reaction
zone and filtered to separate the precipitate from 50
the aqueous solution. The precipitate is assayed and is
found to comprise in excess of 99 percent aluminum
phosphate and less than 0.001 percent U30g. The pH
level of the filtrate is measured and is found to be about
0.2. The filtrate is analyzed and is found to contain 23 55
gil Ah03 and 3.7 gil P20S.
The formation of the aluminum phosphate precipitate
is found to regenerate an amount of mineral acid equivalent
to in excess of 800 lbs. of 96 percent sulfuric acid
per ton of original leached zone material. This repre- 60
sents in excess of about 50 percent of the acid necessary
to solubilize a similar quantity of the leached zone material.
65
A representative sample of an aqueous solution of
reaction products resulting from sulfuric acid leaching
of a minus 150 mesh fraction of calcined leached zone
6
material is introduced into a reaction zone comprising a
modified Parr autoclave. The solution is formed by
contacting 600 lbs. of 96 percent H2S04 with one ton of
leached zone material that is calcined prior to contact
with the acid. The aqueous solution is analyzed and is
found to contain 100 gil Ah03, 120 gil P20S, 0.3 gil
U30g and have a pH of about 1.3. The aqueous solution
is heated in tlIe reaction zone to a temperature of about
200· C. while maintaining the autogenic pressure of the
aqueous solution. The solution is maintained at the elevated
temperature for about 5 minutes to effect precipitation
of crystalline aluminum phosphate in the solution
to form a slurry. The slurry is withdrawn from the
reaction zone and filtered to separate the precipitate
from the aqueous solution. The precipitate is assayed
and is found to comprise in excess of 99 percent aluminum
phosphate and less than 0.001 percent U30g. The
pH level of the filtrate is measured and is found to be
about 0.5. The filtrate is analyzed and is found to contain
39.5 gil Ah03 and 36.4 gil P20S.
The formation ofthe aluminum phosphate precipitate
is found to regenerate an amount of mineral acid equivalent
to in excess of 420 lbs. of 96 percent sulfuric acid
per ton of original leached zone material. This represents
in excess of about 70 percent of the acid necessary
to solubilize a similar quantity of the leached zone material.
While the present invention has been described with
respect to what at present are the preferred embodiments
thereof, it will be understood, of course, that
certain changes, substitutions, modifications and the like
can be made therein without departing from its true
scope as defined in the appended claims.
What is claimed is:
1. A process for the regeneration of mineral acid used
to solubilize phosphate ore comprising:
contacting an ore comprising aluminum, phosphorus
and other values including uranium with a leach
solution comprising a mineral acid selected from
the group consisting of phosphoric acid and sulfuric
acid to solubilize at least a portion thereof and
form a solution of spent mineral acid and solubilized
values in association with any non-solubilized
values, said solubilized values including aluminum,
phosphorus and uranium;
introducing said solution of spent mineral acid and
solubilized values into a reaction zone; and
heating said solution to said reaction zone to a temperature
in excess of 100· C. while maintaining the
pressure level at least equal to the autogenic pressure
of said solution to cause a substantially uranium-
free precipitate of crystalline aluminum phosphate
to form and to regenerate at least a portion of
said spent mineral acid to form regenerated leach
solution containing solubilized uranium values.
2. The process of claim 1 defined further to include
the steps of:
contacting said regenerated leach solution with an
organic extractant to extract at least a portion of
any solubilized uranium values present therein; and
recovering said extracted uranium values from said
organic extractant.
3. The process of claim 1 defined further to include
the steps of:
contacting said solution of spent mineral acid and
solubilized values, prior to heating said solution,
with an organic extractant it} I'lxtract at least l\
4,402,919
~
9. The process of claim 8 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 150° C. to about 200° C.
10. The process of claim 8 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 180° C. to about 200° C.
11. The process of claim 8 wherein at least 50 percent
of the mineral acid employed to solubilize the ore is
regenerated.
12. The process of claim 8 wherein at least 70 percent
of the mineral acid employed to solubilize the ore is
regenerated.
13. The process of claim 8 defined further to include
the steps of:
contacting said aqueous solution of reaction products,
prior to heating in said reaction zone, with an organic
extractant to extract at least a portion of any
solubilized uranium values present in said aqueous
solution; and
recovering said uranium values from said organic
extractant.
14. The process of claim 8 defined further to include
the steps of:
contacting said regenerated aqueous leach solution
with an organic extractant to extract at least a
portion of any solubilized uranium values present
therein; and
recovering said extracted uranium values from said
organic extractant.
15. The process of claim 2 defined further to include
the step of:
contacting fresh ore with the uranium-depleted regenerated
leach solution to solubilize at least a
portion of said fresh ore.
16. The process of claim 3 defined further to include
the step of:
contacting fresh ore with the regenerated leach solution
to solubilize at least a portion ofSaid fresh ore.
17. The process of claim 13 defined further to include
the step of:
contacting fresh ore with the regenerated leach solution
to solubilize at least a portion of said fresh ore.
18. The process of claim 14 defined further to include
the step of:
contacting fresh ore with the uranium-depleted regenerated
leach solution to solubilize at least a
portion of said fresh ore.
* * :(I: * *
10
7
portion of any solubilized uranium values present
in said solution; and
recovering said uranium values from said organic
extractant.
4. The process of claim 1 wherein the ore comprises 5
leached zone material.
5. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 100° C. to about 200° C.
6. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 150° C. to about 200° C.
7. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the 15
range of from about 180° C. to about 200° C.
8. A process for the regeneration of mineral acid used
to solubilize phosphate ore comprising: .
separating an ore comprising aluminum, phosphorus,
uranium and other elements into at least two frac- 20
tions, at least one of said fractions having an average
ore particle of a size capable of passing through
a U.S. Standard 150 mesh screen;
contacting said fraction capable of passage through a 25
U.S. Standard 150 mesh screen with an aqueous
leach solution comprising a mineral acid selected
from the group consisting of phosphoric acid and
sulfuric acid to solubilize at least a portion thereof
and form an aqueous solution of reaction products 30
comprising spent mineral acid and solubilized values,
said solubilized values including aluminum,
phosphorus and uranium;
separating said aqueous solution of reaction products
from any unsolubilized ore to provide a substan- 35
tially solids-free solution of reaction products;
introducing said substantially solids-free solution of
reaction products into a reaction zone; and
heating said solution in said reaction zone to a tem- 40
perature in the range of from about 100° C. to
about200° C. while maintaining the pressure level
at least equal to the autogenic pressure of said solution
to cause a precipitate of substantially uraniumfree
crystalline aluminum phosphate to fonn and to 45
cause at least a portion of said spent mineral acid to
.regenerate and fonn regenerated aqueous leach
solution containing solubilized uranium values.
50
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